To address the severe deformation of gob-side entries in deep inclined coal seams, this study employs an integrated approach combining field monitoring, physical modeling, and numerical simulation using a Ningxia mine as a case study. The research first elucidates the asymmetric failure mechanism driven by the dip angle and deep stress field. Consequently, a novel collaborative control technology '' full anchor cable-single side new double hole cooperative internal pressure relief '' is proposed. Integrated simulations demonstrate that compared to traditional single-hole methods, the bi-portal strategy more effectively mitigates peak stress and prevents the formation of secondary stress concentrations through a superimposed relief effect. Field industrial tests confirm the technology's efficacy: following implementation, the average surrounding rock deformation decreased from 555 mm to 135.5 mm (a 75.6% reduction), the increase in anchor cable readings was significantly reduced, and the stress levels stabilized in the later stages (with a maximum of 164.2 kN). These results validate the long-term stability of the roadway and provide a robust reference for ground control in similar complex deep geological environments.

In extremely thick coal seam mining, residual coal pillars and associated goafs induce asymmetric instability in underlying roadways, leading to severe floor heave disasters. Unlike previous studies that rely on single stress parameters, this study integrates deviatoric stress invariants J2 and J3 with unloading damage mechanics to evaluate roadway stability, where J2 reflects the magnitude of shear stress concentration and J3 identifies the dominant failure mode, distinguishing between compression-dominated and extension-dominated stress states. A theoretical stress model incorporating the combined effects of residual coal pillars and adjacent goafs was developed. Triaxial unloading experiments demonstrated that unloading trajectory, rather than final stress state, controls progressive damage evolution. Numerical simulations of four roadway layouts revealed symmetric Y-shaped high stress zones beneath pillar goaf interfaces, and a coal pillar spanning roadway layout (CPSRL) was proposed with the lowest stress concentration and damage indices. A multilevel support system combining surface shotcrete, deep grouting, and active–passive reinforcement was implemented.

Stability maintenance of deep roadways in thermally damaged rock masses presents complex challenges under asymmetric stress conditions. This study investigates the failure mechanisms and proposes a control strategy using an integrated methodology of laboratory testing, theoretical analysis, and field validation. Triaxial compression tests with Acoustic Emission monitoring indicated that the mechanical degradation of thermally damaged rock is primarily influenced by the propagation of pre-existing microcracks. An analytical model based on the complex variable method quantified the stress field around a rectangular opening, incorporating the effect of principal stress rotation. The analysis revealed a coupled failure mechanism where stress asymmetry governs the inclined X-shaped plastic failure geometry while degraded rock properties determine the extent of the plastic zone. To address this mechanism, a Reinforcement-Anchorage-Confinement (RAC) collaborative support system was developed. Numerical simulations and a field application demonstrated that the RAC system controls fracture propagation and maintains roadway deformation within acceptable operational limits. This research provides a mechanistic framework for roadway stability control in similar geomechanical environments.

Stability control of deep underground excavations under repeated mining disturbances remains a critical geotechnical challenge. Taking a typical extra-thick coal seam as the engineering background, this study reveals a three-stage evolution mechanism of overburden transformation by integrating physical simulation, theoretical modeling, and numerical analysis. A quantitative load transfer model comprising far-field bending subsidence, near-field key blocks, cantilever beam is established to elucidate the progressive load transfer mechanism. Research results demonstrate that the secondary mining stage represents a critical transition where the roof structure shifts from a self-supporting masonry beam to a load-transferring cantilever. Specifically, the key block rotation angle increases from 2.48° to 22.15°, inducing intense stress concentration. Guided by these findings, a stage-adaptive spatio-temporal graded resistance support system with temporally progressive and spatially complementary characteristics was developed to provide active constraint and deep reinforcement. Field implementation validated the system’s effectiveness, reducing rib deformation by 71.8% and roof-to-floor convergence by 53.0%. This work provides a theoretical basis and practical reference for stability control in deep underground engineering subjected to complex dynamic loads.

To address the limitations of existing subsidence control technologies in coal mining, this study systematically investigates the fundamental principles of cut-and-fill mining, the stability mechanism of the filling body, and the influence law of key parameters on mining engineering effects, through a comprehensive research framework integrating theoretical analysis, similar material simulation and numerical simulation. Firstly, the mechanical characteristics of horizontal and diagonal shear failure of gangue pillars are revealed via theoretical derivation. It is clarified that the diagonal stability of the gangue pillar can be guaranteed when its aspect ratio is ≤0.5, and the lateral constraint of metal mesh can effectively enhance its horizontal stability. Secondly, based on a physical model with a size similarity ratio of 1:100, the overburden failure characteristics are obtained: only local cracks appear in the immediate roof and the basic roof presents gentle subsidence after cut-and-fill mining, which directly verifies the effective control effect of this technology on mining-induced overburden movement and surface subsidence. On this basis, multiple sets of orthogonal tests are designed using FLAC3D software (5.0) to analyze the effects of roof cutting width, filling width and coal seam thickness on roof displacement and filling area stress. Combined with grey correlation analysis, it is determined that coal seam thickness is the most critical factor affecting the mining effect, with the correlation coefficients for roof displacement and filling area stress reaching 0.79 and 0.93, respectively. The research shows that the parameter combination of 10 m roof cutting width + 10 m filling width (Group 10-10-X) can achieve the optimal balance between subsidence control efficiency and filling engineering benefit; for working faces with higher requirements for surface subsidence control, the combination of 5 m roof cutting width + 10 m filling width is recommended. The research results clarify the action mechanism of cut-and-fill mining, optimize the key engineering parameters, and provide a solid theoretical basis and technical support for the engineering popularization of this technology and high-precision surface subsidence control.

Coal wall rib spalling, a frequent safety hazard in underground coal mining, is driven by the complex stress environment induced by mining-induced strata movement and dynamic loading from coal cutting equipment. Its severity is significantly exacerbated with increasing mining height, posing major threats to the safe operation of working faces. This study systematically investigates the failure mechanism and key influencing factors of coal wall rib spalling via an integrated framework combining theoretical analysis, FLAC³D finite difference numerical simulation, L25(5⁶) orthogonal testing, and grey relational analysis (GRA). Two dominant failure modes were identified: shear slip failure and tensile spalling failure, and shear slip failure is sensitive to factors such as coal bulk density and cohesion, while tensile spalling failure is more sensitive to factors including mining height. Six key influencing factors were determined and classified into three categories: coal seam occurrence conditions (mining height, coal bulk density), coal mass mechanical properties (cohesion, internal friction angle), and support parameters (hydraulic support intensity, shield guard horizontal supporting stress). Range analysis and GRA yielded highly consistent results: the grey relational degree of the factors is ranked as cohesion > internal friction angle > mining height > coal bulk density > hydraulic support intensity > shield guard horizontal supporting stress. This indicates that coal mass properties exert the dominant control on coal wall stability, followed by seam occurrence conditions, while support parameters have the weakest influence. The findings clarify the priority of rib spalling prevention, and provide critical theoretical guidance for formulating coal wall stability control schemes in high-cut fully mechanized mining faces.

The distribution characteristics of the deviatoric stress tensor (DST) field in the surrounding rock of the end-mining retreat roadway (ERR) are intricate, exerting a substantial influence on the stability of the ERR’s surrounding rock. Taking the fully mechanized working face and its retreat roadway in Wutong Coal Mine as the engineering backdrop, initially, a numerical model of the DST in the end-mining coal pillar (ECP) was established to analyze its evolutionary pattern. Based on the response surface methodology (RSM) model, the significance of each influencing factor was examined. Subsequently, the three dimensional stress expression above the ECP was deduced, and the invariants and distortion energy (DE) of the DST at any position within the ECP were ascertained. Then, the expressions for the cutting height and cutting angle were derived, and a novel time step control technology (TSCT) for the ERR’s surrounding rock was put forward, based on the real-time evolution characteristics of DST, dynamic matching of support strategies and roof cutting parameters is achieved through two stages of passive reinforcement and active roof cutting to achieve collaborative control. Compared with traditional static design or single time step methods, this technology reduces the deformation rate of surrounding rock by about 84% through step-by-step and timely application of support and cutting. Ultimately, according to the field measured data, it was demonstrated that this technology can effectively mitigate the deformation of the ERR’s surrounding rock.

Asymmetric floor heave in gob-side entries of thick coal seams is controlled by rib-to-rib vertical-stress imbalance acting together with in-situ horizontal stress. We propose a layout-first mitigation strategy in which, under a staggered-elevation mining scheme, the gob-side roadway is sited as a negative coal pillar within a low-stress trapezoidal coal mass protected by the hinged semi-arch formed by key block B. A Rankine-based limit-equilibrium model is developed to quantify the active–passive slip mechanism of floor failure and to rank the sensitivity of key parameters. UDEC numerical simulations and 1:100 physical similarity tests are then used to resolve the associated stress redistribution, showing outward migration of peak abutment stress, a widened low-stress belt, and a transition from deep, through-going floor damage to shallow, segmented damage. To enable reproducible comparisons across layouts, three dimensionless indicators (Sf, Rpeak, and D) were introduced, together with a lightweight statistical workflow. Field monitoring on two adjacent panels at the Zhenchengdi Mine further verified the effectiveness of the NCP layout. Compared with the conventional gob-side roadway with a 20 m coal pillar, the NCP layout increased the roadway-to-peak distance to about 12.6 m, widened the rupture-zone width to about 7.8 m, reduced surrounding-rock convergence, and prevented observable floor heave within the monitored advance. These results indicate that the sidewall stress-concentration factors (K, K′) have a stronger influence than the lateral pressure coefficient (λ).

Aiming at the control difficulty of large deformation driven by different mechanisms in the solid coal rib and the coal pillar rib of gob-side entry driving under the influence of superimposed strong mining-induced stresses from opposing mining and excavation in thick coal seams, a comprehensive research approach combining theoretical analysis, numerical simulation, physical similarity simulation, and field industrial tests was adopted. The main conclusions are as follows: ①The mechanisms of large deformation driven by different causes were revealed: the "superimposed mining stress-driven type" for the solid coal rib and the “superimposed overlying strata structural rotation-driven type” for the coal pillar rib. Based on this, a “roof-rib” dual pressure relief collaborative control technology was proposed. ②A comprehensive control scheme was formulated, integrating pre-splitting roof cutting on the gob side, internal pressure relief and anchoring synergy in the solid coal rib, and steel shed-cable coupling in the coal pillar rib. Specifically, the roof cutting parameters include a height of 10m, an elevation angle of 90°, and a cutting hole spacing of 0.6m; the large-diameter pressure relief boreholes feature a starting depth of 6m, a length of 4m, a spacing of 3m, and a diameter of 1m. ③The collaborative control scheme must be implemented no later than 30m before the intersection of the opposing mining and driving faces, covering a total implementation length of 80m within the strong superposition zone. ④Numerical simulations demonstrate that the stress concentration degree of the two ribs is significantly reduced by 44.1% after the adoption of this technology. Field strata behavior monitoring indicates that the convergence of the roadway roof-to-floor and the two ribs is reduced by approximately 74% and 77%, respectively, compared with the original scheme. The research results can provide an important reference for the surrounding rock control of other roadways subjected to strong mining-induced stresses.

Given the problem of the solid coal side produces strong mine pressure in the stage of strong mining and lagging dynamic pressure in the gob-side entry retaining (GER). Using the W3233 ventilation roadway in the West-Third mining area as a case study, the research employs numerical simulation, field observation, and engineering analogy. Key findings include: (1) The optimal plan reduces maximum stress on the coal side and shifts it inward, primarily by severing the main roof and alleviating the gangue wall load, offering limited pressure relief on the solid coal rib. (2) When hole-creating depth is too close to the solid coal rib (b = 6 m), it compromises integrity. If the depth exceeds the stress maximum region (b = 10 m), it fails to effectively sever the peak stress area. However, a depth within the stress maximum region (b = 8 m) effectively absorbs and transfers stress, with a buffer area of approximately 3.7 m from the external support area. (3) The length of the hole (l = 2, 4, 6 m) directly affects deformation absorption and pressure relief, while longitudinal row spacing (s = 2.7, 3.6, 4.5 m) has an inverse effect. For optimal results, a hole length of 4 m and a row spacing of 3.6 m are recommended. (4) After 150 days, the new bag pressure regulating system maintains about 40 % of the absorbable deformation space in the LPRH, confirming simulation results. Observations show that external support forces remain stable at around 250 kN, and coal rib displacement is controlled within 350 mm, over 65 % less than similar GER sections with comparable geological conditions. This research introduces a novel technology and methodology for managing large deformations of the solid coal rib in gob-side dynamic pressure roadways, providing significant reference value.

Given the lack of systematic research on the movement of overburden strata and damage characteristics of floor residual coal in the process of thick coal seam top coal caving, this study takes the Anping Coal Mine as the engineering background. The spatiotemporal evolution of overburden collapse, force chain transmission, and floor stress redistribution was investigated via the FLAC3D‒PFC coupled numerical method. Owing to the large mining space of the thick coal seam, the collapse of the roof has the progressive characteristic of “first two ends, then the middle”, with a measured collapse angle of approximately 57°, forming an evident caving band and residual coal accumulation zones. Moreover, the force chain network in the goaf evolves from sparse to dense, gradually forming a stable load-bearing structure. In addition, the porosity of the floor fluctuates from a wide range (0.15–0.43) to a more stabilized level (0.18–0.40) after compaction. Correspondingly, the vertical stress in the center of the residual coal floor increases from nearly 0 MPa to 1–3 MPa as the overburden load is transferred downward. The research results have important reference value for the destabilization mechanism of surrounding rock in thick coal seam mined-out areas and the design of residual coal remining and provide feasible ideas for improving mine productivity and guaranteeing safe production and the sustainable development of coal resources.

Accurate mining stress monitoring is key for mine layout optimization, roadway support structure design and impact ground pressure prevention and control. However, the plastic and fracture zones generated during the drilling and installation of borehole stress gauges lead to inaccurate or even ineffective results of traditional mining stress monitoring. To address this problem, this study aimed to systematically reveal the failure mechanism of traditional borehole stress gauges for monitoring mining stress in coal seams and to propose a new stress monitoring method based on grout curing pretreatment, including reshaping the integrity of the surrounding rock of the borehole through grout curing and constructing a stable medium for stress transfer. And the new method of monitoring mining stress in coal seam by borehole stress gauge is studied by laboratory test, numerical simulation, and field application. Both laboratory test and numerical simulation verify the feasibility of the new method of mining stress test in improving the monitoring accuracy of borehole stress gauge, and the larger the range of plastic and fracture zone, the more significant the improvement effect of monitoring accuracy. The new method was applied in the field. The monitoring results show that the traditional monitoring method of borehole stress gauge easily failed, and the value was low, while the monitoring accuracy of borehole stress gauge is significantly improved after grouting solidification. The field monitoring results are basically consistent with the laboratory test and numerical simulation results. This method can significantly enhance the measurement accuracy of borehole stress gauges. The results can promote the optimization of mine development layouts, the design of roadway support structures and the prevention and control of rock bursts.

Double-layer island working face main roadway coal pillars are affected by complex mining stress superposition, when different coal pillar width combinations, the surrounding rock stress field will produce different degrees of regional loading increase effect; the study of the surrounding rock stress field regional superposition loading increase law is meaningful to explaining the failure mode of the roadway and determining the critical control area. This study combines numerical simulation with on-site monitoring and other methods and draws the following conclusions: The superimposed loading increase law (“decreasing” → “increasing”) of the abutment pressure and deviatoric stress in the lower coal seam of the double-layer island working face during the mining; the type of the principal stress deflection in the advance working face region; and by obtaining the three types of development morphology of the deviatoric stress peak zone of the roadway and its corresponding nine evolution modes (one type of circular tube → four types of inverse hyperbolic body → four types of hyperbolic body) in the double-layered island working face mining. Indicated the critical reinforcement area corresponding to the main roadway when at different combinations of coal pillar widths; determined the main track roadway protective coal pillars width for 40 m and the shape of the roadway peak deviatoric stress zone is the inverse class hyperbolic body mode; according to the evolution mode of the peak deviatoric stress zone, determined the synergistic failure control program for the asymmetric critical zone of the roadway surrounding rock which is a targeted scientific support method; after the feedback of on-site monitoring and, the support program is reasonable and effective.

Aiming at the problem of large deformation and instability control of narrow coal pillar in the vertical stratification section of extra-thick coal seam, the stability and control technology of the surrounding rock of narrow coal pillar in extra-thick coal seam are studied by combining theoretical analysis, numerical simulation and field test. Based on the calculation of the triaxial stress in the narrow coal pillar, the yield criterion of the plane strain of the narrow coal pillar is obtained based on the Mohr-Coulomb failure strength, the failure characteristics and the evolution law of the damage degree of the narrow coal pillar are qualitatively analyzed, and the instability mechanism and size effect of the narrow coal pillar are further analyzed. Finally, the key technologies of surrounding rock control of the narrow coal pillar are given. It is verified by numerical simulation and field test. The research results show that ① The damage degree of the middle part of the narrow coal pillar is greater than that of the two sides, and the damage degree of the middle part from the top to the bottom develops from serious to slight. With the increase of width-height ratio, the damage degree and the proportion of serious damage area in the narrow coal pillar gradually decrease. When the width-height ratio is greater than 1∶1, a large range of slight damage area begins to appear in the middle and lower part of the narrow coal pillar. When the width-height ratio is greater than 5∶3, the proportion of slight damage area exceeds 50%. When C≥3 MPa, or φ≥20°, the damage degree of the two sides of the narrow coal pillar becomes slight. ② The side of the mining roadway in the working face is a low-stress bearing area of the narrow coal pillar. Under the action of high stress in the upper part, the coal body in the large-scale low-stress bearing area moves to the roadway, resulting in a continuous large deformation of the two sides, which in turn affects the stability of the roof. The height of coal pillar is the main controlling factor for the stability of the narrow coal pillar in the middle and bottom layered sections. The increase of width-height ratio of the narrow coal pillar has little influence on the bearing strength of high strength bearing area, while the bearing strength of low strength bearing area increases obviously. Reasonable width-height ratio of the narrow coal pillar can balance the proportion of high and low strength bearing area in coal pillar. ③ The stability of surrounding rock should be considered in the design of layered narrow coal pillars when both sides of the working face are mined out, so as to ensure that the range of high-strength bearing area inside the coal pillar is greater than half of the width and height of the coal pillar. Through the joint strengthening support measures, multiple joint control areas from shallow to deep are formed in the coal pillar to jointly maintain the self-stability of the coal pillar.

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The section span of the withdrawal space of fully mechanized top coal caving in an extra-thick coal seam is large, and with the gradual withdrawal of the hydraulic support, a series of strong dynamic pressure disasters occur in the withdrawal space, and the difficulty of surrounding rock support control increases sharply. In order to study the control mechanism of surrounding rock in the final mining withdrawal space in detail and put forward a reasonable support technology scheme, taking the large-section withdrawal space of an 8309 fully mechanized caving face in an extra-thick coal seam of a mine as the research object—through the theoretical investigation of whether the key blocks of the main roof are stably hinged under varied stopping coal caving distances and fracture locations of the main roof—the reasonable and optimal stopping coal caving distances and roadway formation time are determined. Using numerical simulation and similar simulation methods, the vertical stress and the maximum shear stress research indicators were introduced to verify the accuracy of the theoretical analysis results. The results show the following: (1) The reasonable stopping coal caving span is 1~2 times the cycle weighting interval, the best stopping coal caving distance in this geological condition is 30 m, and the best fracture position of the main roof is located above the goaf. (2) The migration of overlying strata in the withdrawal space has obvious zoning characteristics, and the zoning is as follows: a stopping coal caving area, support area of the hydraulic support, withdrawal channel area, and stopping coal pillar area. (3) According to the zoning characteristics of overlying strata movement, the asymmetric zoning support control scheme of the withdrawal space is proposed. The field monitoring results show that the maximum roof subsidence in the withdrawal space is 151 mm, the maximum internal squeezing amount of the stopping coal pillar is 82 mm, and the supporting and anchoring effect of each partition in the withdrawal space is good. The set of partition asymmetric support control schemes has been successfully applied to field practice.

Under double-seam mining, the main roadway surrounding rock is affected by the superposition of the advanced stress of the two-seam coal working faces. The stress superposition mode and degree are of great significance to the width calculation of the protective coal pillar and the determination of the critical control direction of the surrounding rock. This paper uses theoretical analysis, numerical simulation, and site engineering practice to carry out targeted research. The conclusions are as follows: Under different lateral pressure coefficients, the superposition evolution law of maximum principal stress direction of two coal seams with different offsets; Two developmental trends and three types of evolution models of J2 peak zone (the critical area of the stress increase and deflection changes) under different superimposed loading modes are summarized. Based on the typical asymmetric evolution model of the J2 peak zone, an asymmetric truss-cable co-anchoring method is proposed aimed at the J2 critical zone. The field monitoring results show that the main roadway surrounding rock is stable after support when the upper coal seam protective coal pillar is left 80 m, and the lower one is 60 m wide. It is of great reference importance for similar engineering practices.

In order to solve the support problem of deep soft crushed coal roadway, a concentrated cavern in a mining station of a mine is taken as the test object. Based on the analysis and summary of the field observation data and the law of rock pressure appearance, a new technology of pressure relief anchoring with the main body of “initiative support + borehole pressure relief” is proposed. This new technology will carry out strong active support in the shallow part of the surrounding rock and excavate a row of low-density large-diameter pressure relief boreholes in the deep coal body of the roadway ribs. The numerical analysis model is established by FLAC3D, and the second invariant of deviatoric stress (J2) is used as the analysis index to elaborate the influence of different borehole parameters on the pressure relief effect of roadway surrounding rock. The results show that different borehole parameters have different effects on roadway pressure relief, that is, borehole depth > borehole length > borehole spacing. After the borehole process is used to relieve the pressure of the surrounding rock, the deformation of the mining roadway side in the subsequent observation process is always controlled within the range of 100 mm, and the shallow surrounding rock support system is effectively protected. The comprehensive control effect is very obvious. Therefore, the field practice proves that the supporting technology can effectively solve the problem of large deformation support of similar roadway surrounding rock.

China has abundant coal resources, and the distribution of coal seams is complex. Thick coal seams account for more than 45% of all coal seams. Fully mechanized top coal caving mining has the advantages of large production, high efficiency, and low cost. In fully mechanized caving mining, especially in fully mechanized caving mining of extra-thick coal seams, the mining space is ample, the mine pressure is severe, and the roadway maintenance is complex. As a result, it is necessary to summarize and discuss the gob-side entry driving of fully mechanized caving in theory and technology, which will help to promote the further development of fully mechanized caving gob-side entry driving technology. First, in recent years, the research hotspots of gob-side entry driving have focused on the deformation mechanism and the control method of the roadway surrounding rock. Secondly, this paper discusses the theoretical models of the “triangle-block” and “beam” for the activity law of the overlying strata in gob-side entry driving, including the lateral breaking “large structure” model, compound key triangle block structure model in the middle and low position, the high and low right angle key block stability mechanics model, elastic foundation beam model, low-level combined cantilever beam + high-level multilayer masonry beam structure model, and the vertical triangular slip zone structure model. It introduces the “internal and external stress field theory” and the “stress limit equilibrium zone model”. Thirdly, it summarizes several numerical simulation analysis methods in different conditions or research focuses and selects appropriate constitutive models and simulation software. Finally, it introduces surrounding rock control technology, including two ribs, the roof, and under challenging conditions. It provides a method reference for support in similar projects.